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By Inkap Choi

Asbestos is a broad term embracing a number of fibrous silicate minerals. These minerals look extremely fragile, yet their fibers have a tensile strength equal to that of piano wire. Chrysotile asbestos is the only mineral that can be woven into cloth, and its fibrous structure is, if anything, even more amazing than its remarkable ability to withstand heat. Because of its unique physical properties asbestos has achieved great industrial importance, and more than 3,000 current uses are known, ranging from use in the kitchen to use in space missiles. Because their minuscule fibers are eminently respirable, asbestos minerals have also found their way into the lungs of humans, where they apparently produce serious biological effects such as asbestosis and lung cancers. 1, 2,3 Because of both their usefulness and their hazardous properties, these minerals have attracted a good deal of attention from research workers during recent years. For example, the surface properties of asbestos minerals have been studied by many investigators. 4, 5, 6, 7, 8, 9 However, even for the important mineral chrysotile, the literature data are more or less conflicting. Pundsack4 measured a PZC (point of zero charge) of 10.1; Martinez and Zucker5 measured 11.8; Chwastiak reports 10.8-11.0; and Raddick measured three sepa-rate apparent PZCls at about pH 8.3, 10.0 and 10.9. More recently Trivedi8 reports 10.11. Naumann and Dresher9 have studied surface charge enhancement of chrysotile by metal ions, and describe how the mineral reacts with contaminantions. In the present paper, the mode of dissolution of magnesium and hydroxyl ion from asbestos minerals placed in water was studied by kinetic experiments involving rate at which these items leave the minerals. Also, some study was made of interaction between dissolved species.

Jan 1, 1972

By Dewa Putu Agus Sudarmaja, Widi Astuti, Januarti Jaya Eka Putri, Yayat Iman Supriyatna, Himawan Tri Bayu Murti Petrus, Siti Nurul Aisyiyah Jenie, Ferian Anggara, Anisa Maulidia, Vincent Sutresno Hadi Sujoto, Sutijan, Wayan Warmada

With the rapid development of renewable energy technology, such as electricity reserves to store energy for electric and battery-electric vehicles, environmentally friendly energy sources are expected to continue to be tapped to meet mobility and stationary energy needs. Lithium is the most attractive candidate for energy storage development. The presence of lithium in the Sidoarjo mud in Indonesia makes it interesting for use as an alternative source of raw material. In this research work, sulfuric acid was used as a solvent, and hydrogen peroxide as an oxidation agent. Lithium recovery was studied with a leaching process at various temperatures (30, 45 and 60 °C) and sulfuric acid concentrations (0.5, 1, 2, 3, 4 and 6 mol/L). Increases in temperature, sulfuric acid concentration and hydrogen peroxide concentration were found to directly affect the lithium recovery. An increase in sulfuric acid concentration is directly proportional to the recovery of lithium in the solution. Optimum conditions were obtained at a sulfuric acid concentration of 6 mol/L and temperature of 60 °C. The shrinking core model was used to study leaching kinetics, and ash layer diffusion acted as a rate-controlling process. The activation energy of the leaching process was found to be 38.90 kJ/mol.

Jun 1, 2024

By Dewa Putu Agus Sudarmaja, Widi Astuti, Januarti Jaya Eka Putri, Yayat Iman Supriyatna, Himawan Tri Bayu Murti Petrus, I Wayan Warmada, Siti Nurul Aisyiyah Jenie, Ferian Anggara, Anisa Maulidia, Vincent Sutresno Hadi Sujoto, Sutijan

With the rapid development of renewable energy technology, such as electricity reserves, to store energy for electric or electric vehicles, environmentally friendly energy sources are expected to continue to be fulfilled according to mobility and stationary energy needs. Lithium is the most attractive candidate for energy storage development. The presence of lithium in Indonesia in the Sidoarjo mud content is interesting to be used as an alternative source of raw materials. During the research, sulfuric acid acted as a solvent, and hydrogen peroxide acted as an oxidation agent. Lithium recovery will be studied using a leaching process at various temperatures (30, 45, and 60 °C) and sulfuric acid concentrations (0.5, 1, 2, 3, 4, and 6 mol/L). The increasing temperature, sulfuric acid concentration, and hydrogen peroxide concentration will directly affect the lithium recovery. An increase in sulfuric acid concentration will be directly proportional to the recovery of lithium in the solution. Optimum conditions were obtained at varying 6 mol/L sulfuric acid concentrations and 60 °C temperature variations. The shrinking core model (SCM) studied leaching kinetics, and ash layer diffusion acted as a rate-controlling process. The activation energy of the leaching process was obtained at 38.90 kJ/mol.

Jul 25, 2023

By R. Baker

A novel electrochemical half-cell and corresponding working electrode were designed to evaluate the electro-catalytic activities of catalysts in an environment close to that of an operating proton exchange or polymer electrolyte fuel cell (PEMFC). Cyclic voltammetry (CV) and electrochemical impedance spectroscopy (EIS) techniques were employed using the novel cell to characterize various methanol electro-oxidation parameters that contribute to the overall reaction on a fuel cell anode. The effect of temperature on methanol oxidation kinetics was examined in the potential range of 0.31 to 0.65V (vs. NHE). The data was collected in the temperature range of 20 to 80°C. Based on the experimental analysis and literature data, an equivalent circuit model (ECM) for methanol oxidation was proposed and was used to fit the EIS experimental data. The experimental data and ECM electrochemical data was used to extract electrochemical kinetic parameters for methanol oxidation such as the exchange current density, Tafel slope, and energy of activation for the reaction. Kinetic data, especially at different temperatures, is very limited in the literature.

Jan 1, 2005

By T. T. Chiweshe, L. S. Mokhahlane, M. Welman-Purchase

This study examined an alternative technique of extracting tungsten from wolframite ore using a fusion method with ammonium phosphate salt. Current techniques use alkali salts to decompose tungsten ores and use various purification methods to produce ammonium paratungstate, which is a key intermediate precursor in the extraction of tungsten. In this study, a two-step extraction process, which involves thermal decomposition of the ore using fusion and a separation stage, is examined. The novelty of this technique is the ability to completely decompose wolframite ore and form an unstable tungsten oxide phosphate (WO2(PO3)2) compound. The instability of WO2(PO3)2 in deionized water enables selective isolation of tungsten as insoluble precipitates of WO3 and WO2. The decomposition reaction of WO2(PO3)2 is influenced by the presence of oxygen and determined to be first order under ambient conditions. Scanning electron micrographs and X-ray diffraction analysis showed above 80% of the tungsten precipitated compared with the amount that remained in the filtrate solution.

Jul 10, 2024

By Gerald Onyedika, Tochi Onwumere, Martin Ogwuegbu

"The kinetic study for the leaching of iron from cassiterite ore in KOH solution was studied. The effects of KOH concentration, particle size and reaction temperature on the leaching rate of iron were also examined. The results of leaching studies showed that the leaching of iron from cassiterite ore in KOH solution is significantly influenced by the reaction temperature, KOH concentration, and decrease in particle size. The study showed that 90.62% of iron ore was leached in 8M KOH solution at 80°C within 120 minutes. The experimental data were well evaluated using the shrinking model for surface chemical reaction and diffusion control through the product layer. The activation energies for the iron ore dissolution using the two models were calculated to be 32.06KJmor1 and 97.97 KJmor' respectively.IntroductionIn recent times, there has been increasing attention on the kinetic study of mineral dissolution. In most of these studies, the experimental results were analyzed with a suitable shrinking core model under the assumption that the ore is a homogeneous spherical solid phase [1-4]. A study of the dissolution kinetics of a Nigerian cassiterite ore by hydrochloric acid showed that increase in the hydrochloric acid concentration is accompanied with increase in the amount of cassiterite ore dissolved at various times [4]. The optimal concentration was found to be 4.0M HCl, after which there was a decline in the amount of the ore dissolved, which showed that the concentration of the reagent has a significant effect on the leaching rate of the ore. Also in studying the effect of stirring speed on the rate of cassiterite ore dissolution while keeping other conditions constant for 120 minutes, the result showed that the rate of cassiterite ore dissolution was dependent on the stirring speed over a range of 0 to 630 rpm. Above 360 rpm, the stirring speed no longer had any effect on the solid dissolution. The dissolution rates of cassiterite were analyzed using a suitable shrinking model. The results showed that the reaction was controlled by the diffusion of tin through the porous product layer, a reaction order with respect to hydrogen ion concentration of 0.92 M and activation energy of 50.05KJmor'. In another kinetic study on the leaching of a low grade niobium-tantalum ore with concentrated KOH solution the effect ofreaction temperature, KOH concentration, agitation speed, particle size and alkali-to-ore mass ratio on the dissolution rate of niobium were examined [5]. The result of niobium leaching from low grade niobium-tantalum ore with concentrated KOH, solution at different agitation speeds showed that the dissolution rate increased with increase in agitation rate. In the case of particle size effect, the results show that small particles require much shorter time for complete dissolution than the large ones."

Jan 1, 2013

By J. B. Hiskey

The principle oxide minerals of molybdenum are wulfenite (PbMoO4), ferrimolybdite (Fe2(MOO 4)3.nH2O), powellite (CaMoO4), and ilsemannite (Mo304). Wulfenite is found in Arizona, New Mexico, Nevada, Utah, California, and Montana. It is also found in the states of Sonora and Chihuahua, Mexico. Prior to 1915 wulfenite ores were the primary source of molybdenum. The solubility of wulfenite in hot sodium sulfide was first noted by Bonardi(1) during attempts to sulfidize wulfenite material prior to flotation. He identified lead sulfide and soluble sodium molybdate as reaction products. Bacon (2) sulfidized wulfenite in preparation for flotation and by X-ray analysis found the reaction product to be lead sulfide. Holmes(3) made an extensive investigation of the extraction of molybdenum from wulfenite concentrates by leaching. Nearly 100% molybdenum extraction was possible from -100 mesh samples leached for 15 min with 10% sodium sulfide solution at room temperature. Molybdenum extraction from -100 mesh samples after leaching 30 min at a temperature of 45°C using 2, 4, and 6% sodium sulfide was about 32, 79, and 98% respectively. The extraction showed little sensitivity to time, temperature, or particle size using 10" sodium sulfide solution.

Jan 1, 1976

By A. P. Mehta

The role of microorganisms in the leaching of metal sulfides is well established (Torma, 1977; Murr, 1980). However, this established knowledge applies mostly to chemolithotropic microorganisms such as Thiobacillus ferrooxidans which function optimally in leaching catalysis between 25-35° C, and in a pH range of 2-2.5. Although leaching with thermophilic microorganisms capable of extending the range of catalytic viability beyond 35° C is also documented, there are few if any attempts to optimize or extend sulfide concentrate leaching with thermophilic microorganisms. When mixed metal sulfides or other materials having different rest potentials (referenced to the standard hydrogen electrode, SHE, in the electromotive series) make contact in a conducting, aqueous medium, the contacting members will act as an electrochemical cell, with the member having the higher rest potential assuming the role of an electrical cathode. In leaching reactions involving such galvanic couples, the cathode becomes passivated, while the anodic member reacts vigorously due to the electro- chemical reaction which occurs (Hiskey and Wads- worth, 1975). Serry, eta1 (1978) observed that when chalcopyrite and pyrite were in intimate contact in natural ores, the chalcopyrite reacted vigorously by galvanic conversion as a result of the higher rest potential (~ 0.63 V) as compared to the chalcopyrite (~ 0.53 V). In addition, it was postulated that because of the elemental sulfur which can form on the anodic chalcopyrite in this galvanic conversion process, reactions in acid-bacterial medium would be expected to have accelerated copper extraction because of the conversion of any sulfur product layer on the chalcopyrite anode to sulfuric acid according to the following general reactions: [MS?M2+ + S° + 2e- (1) and 2S° + 302 + 2H20 bacteria?-2H2SO 4 (2)] where M in reaction (1) designates the metal (or metals in the specific case of chalcopyrite) to be released to solution by galvanic conversion.

Jan 1, 1982

By S. N. Shahcheraghi

Bioleaching method makes use of bacteria in solving metallic sulphides, so that these bacteria tend to convert insoluble metal sulphides into soluble sulphates. Since the kinetics description of this process is inevitable for better interpretation of relation between metabolism and growth of live organisms, as well as optimization of metals extraction processes, in this paper, this description is represented and designed in mathematic (kinetics) model for copper sulphide minerals based on Monod kinetics model. According to this model, special growth rate of bacteria (?) is a hyperbola function of restrictive substrate concentration (S). By Solving and developing of Monod kinetics model, a comprehensive kinetics model for bioleaching of copper sulphide minerals results. The designed is also applicable for other metal sulphides.

Sep 1, 2012

By José Roberto de Oliveira

In a smelting reduction process the reduction reaction occurs predominantly in the liquid state. The main reduction agent is the carbon from the carbon saturated iron bath. The aim of the present work was to determine the velocity equation and the rate constant of the reduction rate of MnO content in a Mn0-SiO2-Fe0 slag by iron saturated carbon and the carbon dissolution rate in a Fe-Mn bath. The manganese content in the iron bath was in the range of 5 to 15 wt.%. It was observed that the rates of carbon dissolution in all Fe-Mn baths are faster than the rates of MnO reduction by the carbon saturated liquid iron.

Jan 1, 1999

By M. Zhang, H. Ni, F. Zheng, M. Guo

"In the present work, we investigate the effects of roasting temperature and time on the transformation of the titanium (Ti)-bearing phase and magnesium aluminate spinel (MgAl2O4) phase in Ti-bearing electric arc furnace molten slag during the molten sodium hydroxide (NaOH) treatment process. With applied temperature of 700 °C for 60 min, and a mass ratio of Ti-bearing electric arc furnace slag to NaOH, or Rslag/NaOH, of 1:1.2, the conversion ratios of the Ti-bearing phase and the MgAl2O4 phase reached about 98.3 percent and 88.1 percent, respectively. Moreover, investigation of the reaction kinetics indicate that the conversion of these phases during the molten NaOH treatment process is a typical solid-liquid reaction, and it can be theoretically described by the unreacted shrinking core model. The conversion ratio of the Ti-bearing phase was found to be controlled by the chemical reaction step, with an apparent activation energy of 41.76 kJ/mol. The decomposition of the MgAl2O4 phase was found to be controlled by the surface chemical reaction in the early stage of 0 to 20 min from start of reaction, changing to diffusion control in the later stage of 20 to 60 min. The apparent activation energies of the process were determined as 55.88 and 22.72 kJ/mol, respectively.IntroductionChina has rich mineral resources of vanadium–titanium magnetite ore deposits, and the iron and steel industry in the country produces millions of tons of titanium (Ti)-bearing slag every year. The problem of how to dispose of the Ti-bearing slag has attracted much attention, and various approaches have been developed, including the acid process (Li et al., 2007; Zhang et al., 2011; Zhang and Nicol, 2010; Li et al., 2011), alkali fusion process (Han et al., 2012; Middlemas, Fang and Fan, 2013; Nayl, Awwad and Aly, 2009; Lasheen, 2008; Zhang, Qi and Zhang, 2009), carbonization and chlorination process (Liu et al., 2008; Tao et al., 2012; Fu et al., 2012; Liu, 2009), selective crystallization process (Lei and Xue, 2008; Li et al., 2011; Zhang et al., 2007a, 2007b). However, these processes have been beset by various functional, economical and ecological problems (Yu and Chen, 2010). With environmental protection regulations becoming increasingly rigorous, and demand growing for the comprehensive utilization of resources, the focus is on methods to treat Ti-bearing slag with high efficiency and mild reaction conditions (Nazari and Riahi, 2011; Wu et al., 2013)."

Jan 1, 2017

By P. C. Beuria

Iron ore concentrate contains different mineral phases viz. hematite, magnetite, goethite, kaolinite, and gibbsite. Presence of some of these phases like goethite, kaolinite, and gibbsite contribute to the loss on ignition (LOI) content of the sample. Iron ore fines (pellet feed) are pelletized for utilization in DRI & BF processes. In the present pelletization technology i.e., straight grate and grate kiln processes, the LOI in the ore should be within the range of 2.0-2.5%. LOI content in the iron ore concentrate is found to be more than 4% after beneficiation of low-grade iron ore. It is therefore difficult to produce good quality pellet because of high LOI content. The LOI content of the sample is a detrimental factor in the pellet making. For effective pelletization, the LOI content in iron ore fines needs to be reduced before pelletization. It can be reduced by fluidized bed roasting. The objective of present work is to understand the kinetics of removal of LOI from iron ore fines. The effects of bed depth, temperature, and residence time on reduction of LOI content in iron ore fines were studied using muffle furnace. The process of roasting was carried out at four different temperatures varying from 400 to 950 °C with a variable residence time of 30, 45, 60 & 75 minutes and at bed depths of 16, 32, 48 & 64 mm. In all instance of experiments carried out, one of the parameters was kept constant. It was observed that at 850 °C and 30 minutes of residence time, there was sufficient removal of LOI in static bed condition even at a high bed depth of 64 mm. Keywords: LOI, pellet feed, fluidized bed roasting, goethite

Sep 1, 2012

By Zoran S. Markovic

The extensive laboratory investigations were carried out in order to determinate behavior of sulphide and nonsulphide copper bearing particles during the flotation. The present ore sample were contained 0.55% Cu with 20.1% of nonsulphide copper in it and gangue material with higher contents of alumina about 19.07%. In this paper will be presented some results such as: effect of particle size, pH and applied collectors. The main goal was in application of collectors S-7518, as well as mixture of KEX and Z-96. Z-96 is a product from Chemical Industry Zupa Krusevac which was used as auxiliary collector in mixture with xanthates. The best results were done with this mixture in ratio KEX / Z-96 as 77/23 % in area of oxidized minerals flotation, but the collector S-7518 was achieved better results in sulphide mineral flotation. The S-7518 has better selectivity between copper minerals and pyrite comparing to mixture of KEX/Z-96 in mentioned optimal ratio. Flotation recovery and copper grade in concentrate were studied by flotation kinetics tests.

Sep 1, 2012

By S. Itoh, T. Nagasaka, J. Ono

With respect to the recycling of used lead-batteries, a pure lead is required to prepare a lead-calcium alloy electrode for a new maintenance-free battery. The pure lead can be obtained by means of electrolytic and/or pyrometallurgical refining. In the lead electrorefming, antimony is concentrated in anode slime. From the viewpoint of the recycling and recovery of metal values from used lead-batteries, especially from anode slime, oxidation experiments were conducted for pure liquid antimony in the temperature range between 973 and 1373 K to elucidate the oxidation kinetics and mechanism. Furthermore, since lead anode slime generally consists of antimony, lead and bismuth, experiments have been also carried out using antimony-lead-bismuth alloy. The equation relating to the mass transfer characteristics in the gas phase was then applied to analyze the oxidation kinetics. As a result, it was found that the anode slime treatment, that is, an oxidation of antimony concentrated in anode slime followed by vaporization as antimony oxide Sb406, should be operated with higher oxygen partial pressure and sufficient amount of gas flow rate at lower temperature around 1073 K. The feasibility of the oxidation treatment for anode slime has been also discussed.

Jan 1, 2004

Kinetic Study on the Leaching of Vanadium from LD Converter Slag Using Sulfuric Acid In order to reduce the environmental impact due to land disposal of LD (Linz-Donawitz) converter slag from steelmaking plant and to valorise this waste material, an acid leaching study was carried out to assess the effect of several parameters on the kinetics of vanadium dissolution. The characterisation and morphology of the feed investigated using XRD, XRF. The factors influencing leaching vanadium such as particle size, temperature, slurry density and the concentration of sulfuric acid were investigated and optimised. The maximum vanadium recovery about 95 per cent achieved at 70¦C, S/L: 1/15, acid concentration: 3M and for 150 minutes of leaching time. The dissolution of vanadium in sulfuric acid shows that there are two stages in the kinetics of leaching. At the first 15 minutes, there is a sharp increase in the amount of vanadium extracted and at longer leaching time, the extraction increase slightly. In the present work, the shrinking core model (SCM) was used to describe the acid leaching of vanadium compounds. A summary of the controlling regimes in the SCM are examined below for short time (less than 12 min.) and long time (more than 30 min.) stages. So, it was found that the kinetics of leaching in both stages is controlled by solid product diffusion regime.

Sep 13, 2010

By M. C. Ruiz

The conventional pyrometallurgical methods to treat copper sulfide concentrates include generally, one or more oxidizing steps where the formation of SO2 gas can not be prevented. From the economical point of view and the nature of some of the processes in which SO2 is produced, its complete capture is not feasible; therefore an important fraction is released to the atmosphere as fugitive emissions. In this research, the reduction of chalcocite was studied in the presence of lime as an alternative method, which overcomes the pollution problems related to SO2 emissions. The effect of temperature, time, concentration of carbon and CaO on the reduction rate were studied. Among these variables, temperature affects most the reduction rate. Conversions over 95 % can be obtained in less that 20 minutes at temperatures over 1000 ° C. The measured partial pressures of the gaseous components indicated that the chalcocite reduction proceeded through the gaseous intermediate species CO, C02, and the overall kinetics was controlled by the Boudouard reaction. The kinetic model In(1-X)= kt fits well the experimental data. Apparent activation energy of 3 14 kJ/mol was calculated for the temperature range of 900 0C-1050 °C.

Jan 1, 1999

By Bern Klein, Richard W. Lawrence, Scott Frostad

Laboratory and field scale kinetic testing was conducted to evaluate scaling up of laboratory data. Two 20- tonne (field) cells were monitored for internal temperature, pH, conductivity, volume and rate of flow, and water quality. On a surface are. basis, results showed sulfide oxidation rates obtained using 50-kg cells are more sensitive to changes in aeration than those from 1-kg cells. Correcting the laboratory oxidation rates for field temperatures with a pH 7 activation energy, which is in the field drainage pH range, returned oxidation rates an order of magnitude lower than the field cell rates.

Jan 1, 2000

By K. Wongpreedee

Recovery and recrystallization of Au wire can degrade strength and alter conductivity properties during exposure to elevated temperature. Au Deformation Processed Metal Metal Composites (Au DMMC's) are being developed for electronic applications requiring high conductivity and high strength. This paper discusses the relationships between microstructure, strength, and resistivity of Au DMMC's. It also presents the results of a recent study on the elevated temperature stability of Au DMMC's of several compositions. Au-7 vol %Ag, Au-14 vol %Ag, and Au-vol 7%Pt DMMC samples were prepared by a powder metallurgy technique. All specimens were processed into wire by extrusion, swaging, and drawing. The smallest final diameter of these wires was 120 ?m. Thermal stability was analyzed by electrical resistivity measurements to determine the relation between transformation rates and time at elevated temperature. The extensive deformation used to produce these composites reshaped the initially equi-axed powder into filaments that are 30 to 100 nm in diameter and 16 to 180 mm in length. The small filament diameter confers high strength on the composite. The high conductivity can be explained by electrons flowing parallel to the filamentary microstructure aligned with the wire axis. Using the Avrami and Arrhenius equation relationships, it is possible to predict the times and temperatures needed to achieve various degrees of homogenization. The Au DMMC's were found to have good thermal stability compared to conventional coldworked Au interconnection wires. Although these composites will revert to solid solutions if exposed to high temperatures for prolonged time periods, their relative stability is sufficient to allow them to maintain their twophase microstructure during the anticipated lifetime temperature profiles of many products.

Oct 1, 2003

By A. Vidyadhar

Printed circuit boards (PCBs) constitute a majorpart of electrical and electronic equipment containing valuable metals such as Cu, Ni, Au, Ag, Pd, Fe, Sn, Pb, etc. The concentration of base and precious metals in PCBs is observed to be several times more than those in their respective ores. Therefore, the recycling of PCBs is necessary to recover the valuable materials. However, the ultrafine particles pose a big challenge. Removal of them is suggested prior to processing by flotation. Froth flotation is observed to be a promising technique for rejecting the plastics from the comminution product. In the present work, enrichment of ground 1.0 mm PCB powder was investigated through flotation route by varying the operating variables such as frother dosage, pulp density, air flow rate and rotational speed of the impeller. The liberation studies indicate that excellent liberation of metal values is achieved from the non-metallic constituents at -1.0 mm size and below. The particulate system is quite rich in metal value with about 23% total metal content. The non-metallic constituents such as plastics are observed to possess strong hydrophobicity while the metal particles in the pulverized mass are hydrophilic in nature. Froth flotation kinetics is studied in depth with a view to facilitate high rejection of plastics and identify optimum operating conditions for the same. A number of experiments are performed to establish the influence of the operating variables on flotation performance. The conditions for achieving higher yield of the metal-rich fraction with respect to a specified grade are discussed in the light of the experimental results. It is found that the requirement of reagents is negligible which could be a very important factor from the commercial stand point. Single-stage flotation increases the metal content from 23% to over 32% with a mass yield of around 75% and over 90% recovery of metal values. 32% of the materials in the feed could be effectively rejected in the float fraction losing less than 4% metal values. The dependence of kinetics on the process variables is also discussed. It was concluded that a high rotor speed helps in efficiently rejecting the plastics. Frother should be added to help stabilize the froth and enhance the kinetics. A moderate air flow is required while pulp density must be kept low for efficient pre-concentration. It is established that the entire -1.0 mm comminution product could be treated by flotation for generating a pre-concentrate. Keywords: printed circuit boards, recycling, froth flotation, kinetics, recovery

Sep 1, 2012

By Aiyuan Ma, Tu Hu, Chenyu Sun, Yonggang Zuo, Hongying Xia, Jinhui Peng, Libo Zhang, Yongguang Luo

"In this study, the adsorption characteristics of tannin acid in aqueous solution were investigated using regenerated activated carbon (AC), with the aim to remove the tannin acid organic matter from the solution after zinc hydrometallurgy and Ge precipitation. Results showed the temperature has little effect on the adsorption ability (q.) of the regenerated activated carbon. The experimental equilibrium data were tested for the Langmuir and Freundlich isotherms. The adsorption equilibrium was better described by the Langmuir isotherm (R2 > 0.99). Pseudo first- and second-order kinetic models were used for describing kinetic data. It was determined that the kinetics of tannin acid removal was well fitted by second-order reaction.IntroductionGermanium recovery is a complex processes [1]. Hydrometallurgical processes are favoured and these are based on the precipitation of germanium as hydroxide or sulphide or alternately with tannic acid [2]. However, in the germanium recovery process, too much tannin is consumed by the germanium-tannins complexation process, and results in a high content of tannic acid remaining in solution, which is harmful to the valuable metal recovery [3]."

Jan 1, 2014